Processes are known in the prior art which provide for the separation of minerals from mineral-bearing ores.
For example, in known processes used for the separation of copper from copper-bearing ores, illustrated diagrammatically in FIG. 1, non-oxidized ores 20 (which might contain as little as 0.5% copper, and typically contain iron sulfides) are processed in a crusher 22, with water 24, to form a slurry 26. The slurry 26 is then transferred to a flotation cell 28, and subjected to physical action, specifically, air sparging and mixing. As a result of the physical action, a substantial portion of the copper value in the slurry 26 rises to the surface of the flotation cell 28 as a froth 30, and is skimmed therefrom by a paddle mechanism 32, while the waste rock 33 (“gangue”) remains in the bulk, and is ultimately passed from the cell 28 to a dryer 34 and discharged as tailings 36. This process of “froth separation” results from differences in wettability of copper as compared to other minerals, and is typically aided by chemical frothing and collector agents 38 added to the slurry 26, such that the froth 30 from such flotation contains 27 to 36% copper. Methylisobutyl carbonal (MIBC) is a typical frothing agent, and sodium xanthate, fuel oil, and VS M8 (a proprietary formulation) are typical collector agents.
The froth 30 is then fed to an oxygen smelter 40, and the copper and iron sulfides are oxidized at high temperature resulting in impure molten metal 42 (97-99%, copper, with significant amounts of iron oxide) and gaseous sulfur dioxide 44. The impure metal 42 is then transferred to an electrolytic purification unit 46, which separates the impure metal 42 into 99.99% purity copper material 48 and slag 50.
The gaseous sulfur dioxide 44 is collected in a reactor 52 wherein it is scrubber and mixed with water 24 to form sulphuric acid 54. The sulphuric acid 54 is suitably blended with water 24 and used to leach oxidized ores, typically by “heap leaching” an ore pile 56. The resultant copper-bearing acid 58 is known as “pregnant leach solution”. Pregnant leach solution 58 is also obtained by mixing solutions of sulphuric acid 54, in vats 60, with the tailings 36 discharged from flotation operations, to dissolve the trace amounts of copper remaining therein.
The copper is “extracted” from the pregnant leachate 58 by mixing therewith, in a primary extraction step 62, organic solvent 64 (often kerosene) in which copper metal preferentially dissolves. Organic chemical chelators 66, which bind solubilized copper but not impurity metals, such as iron, are also often provided with the organic solvent, to further drive the migration of copper. Hydroxyoximes are exemplary in this regard.
In the primary extraction step 62, the copper is preferentially extracted into the organic phase according to the formula:[CuSO4]aqueous+[2 HR]organic→[CuR2]organic+[H2SO4]aqueous                where HR=copper extractant (chelator)        
The mixed phases are permitted to separate, into a copper-laden organic solvent 68 and a depleted leachate 70.
The depleted leachate 70 is then contacted with additional organic solvent 72 in a secondary extraction step 74, in the manner previously discussed, and allowed to settle, whereupon the phases separate into a lightly-loaded organic (which is recycled as solvent 64 in the primary extraction step) and a barren leachate or raffinate 76.
The barren leachate 76 is delivered to a coalescer 78 to remove therefrom entrained organics 80, which are recycled into the system; the thus-conditioned leachate 82 is then suitable for recycling into the leaching system.
The pregnant organic mixture 68 (produced in the primary extraction step 62) is stripped of its copper in a stripping operation 84 by the addition of an aqueous stripping solution of higher acidity 86 (to reverse the previous equation); after phase separation, a loaded electrolytic solution 88 (“rich electrolyte”) remains, as well as an organic solvent, the latter being recycled as solvent 72 in the secondary extraction 74.
The rich electrolyte 88 is directed to an electrowinning unit 90. Electrowinning consists of the plating of solubilized copper onto the cathode and the evolution of oxygen at the anode. The chemical reactions involved with these processes are shown belowCathode: CuSO4+2 e1−→Cu+SO42−Anode: H2O→2H++0.5 O2+2 e1−
This process results in copper metal 92, and a lean (copper-poor) electrolyte, which is recycled as stripping solution 86.
The combination of leaching, combined with extraction and electrowinning, is commonly known in the art as solvent extraction electrowinning, hereinafter referred to in this specification and in the claims as “SXEW”.
In a known application of the described SXEW process, in both the primary 62 and secondary 74 extraction steps, the combined organic and aqueous phases are delivered through a series of mixing vessels (primary P, second S and tertiary T), and then to a settling tank ST, the primary mixing vessel P being about 8 feet in diameter and 12 feet in height, and stirred by a rotary mixer driven by a 20 horsepower motor, and each of the secondary S and tertiary T mixing vessels being about 12 feet in diameter and height, and stirred by a rotary mixer driven by a 7.5 horsepower motor. (The system of primary P, secondary S and tertiary T mixers, and settling tank ST, is replicated to meet volume flow requirements, with each system processing about 10,000 gpm). This provides a mixing regime wherein the organic and aqueous phases are intimately mixed for a period of time sufficient to allow copper exchange (to maximize copper recovery), yet relatively quickly separate substantially into organic and aqueous phases.
In a known application of the froth flotation process, a plurality of flotation cells 28, each being approximately 5 feet square and 4 feet high, are utilized, with pairs of cells sharing a 50 horsepower motor driving respecting rotary mixers (not shown). This provides a mixing regime sufficient to allow the air bubbles to carry the copper value to the surface.
Various modifications can be made to the rotary mixers in the extractors and in the flotation tanks of the foregoing process. However, the general configurations noted above have been found to provide relatively economical results, and significant variations therefrom can impact adversely upon economies.
For example, an attempt to reduce energy costs by scaling-down the motors for the mixers would have consequent impacts either upon the copper recovery efficiency, or upon available process throughputs.
Specifically, the relatively large motors employed are required to drive the sturdy (and therefore heavy) rotary mixers and shafts that are needed to withstand the torques caused by rotation; lower power motors would demand either lower blade speed or smaller blades, with consequent impacts upon mixing and transfer efficiency.